Rotary kiln reduction of iron oxides with pneumatic feeding of a portion of the charge

ABSTRACT

IMPROVEMENTS OF AN ORE TREATING PROCESS, WHEREIN THE ORE AND THE TREATING AGENT ORE BOTH FED TO A ROTARY KILN WHICH OPERATE AT A TEMPERATURE AND PRESSURE SUFFICIENT TO CARRY OUT SUCH TREATMENT, WHEREIN THE ORE IS FED TO THE TREATING ZONE IN SUCH A MANNER THAT ALL OF IT IS FED AT A GIVEN POINT AND THAT FEEDING IS CARRIED OUT UNDER SUCH CONDITIONS THAT THE ORE BEING FED IS DISTRIBUTED THROUGHOUT AT LEAST A PORTION OF THE ROTARY KILN WITH THE LARGER PARTICLES OF ORE BEING DEPOSITED THE FURTHEST DISTANCE FROM THE DISCHARGE END OF THE KILN AND THE SMALLER PARTICLES, OF ORE BEING DEPOSITED MOST PROXIMATE TO THE DISCHARGE END OF THE KILN SO THAT THE DEPOSITION POINT OF THE ORE PARTICLES DISTANCE FROM THE DISCHARGE END IS PROPORTIONAL TO THE SIZE OF THE PARTICLES BEING DEPOSITED, WHEREBY PROVIDING A RESIDENCE TIME OF THE PARTICLES IN THE KILN PROPORTIONAL TO THE SIZE OF THE PARTICLES.

"United States Patent Office 3,667,933 Patented June 6, 1972 US. Cl. 75-33 3 Claims ABSTRACT OF THE DISCLOSURE Improvements of an ore treating process, wherein the ore and the treating agent are both fed to a rotary kiln which operates at a temperature and pressure suflicient to carry out such treatment, wherein the ore is fed to the treating zone in such a manner that all of it is fed at a given point and that the feeding is carried out under such conditions that the ore being fed is distributed throughout at least a portion of the rotary kiln with the larger particles of ore being deposited the furthest distance from the discharge end of the kiln and the smaller particles, of ore being deposited most proximate to the discharge end of the kiln so that the deposition point of the ore particles distance from. the discharge end is proportional to the size of the particles being deposiited, whereby providing a residence time of the particles in the kiln proportional to the size of the particles.

This application is a continuation of application Ser. No. 559,065, filed June 1, 1966, and now abandoned.

This invention relates to the carrying out of reactions which require heat to be input in order to maintain the reaction. It more particularly refers to reactions in which metal ore is reduced from its combined form to its metallic form. k

Rotary kilns are used to a great etxent in carrying outmetal ore reduction processes as well as other similar reactions. In the usual mode of operations particulate solids to be treated are fed to the rotary kiln together with the treating gas or other material which is used for treating purposes. In prior processes, little or no attention has been paid to the form, shape or size of the ore or other material being treated. Generally, solid particle ores are charged to the kiln in whatever size and size distribution they are available.

It is known that, where a wide particle size distribution feed is used, the kiln operates at a lower efiiciency than is desired. It has been discovered that this is largely due to a combination of facts. The large particle size ore tends to move about the kiln at a relatively fast rate and thus to have a relatively low residence time. Unfortunately, the larger particles often require the greatest or longest treatment in the kiln; but because of their physical size are subjected to the shortest relative treating times.

In an analogous manner, it has been discovered, that the fine particle size ore tends to move about the kiln at a relatively slow rate and thus has a relatively long residence time. Again unfortunately, the smaller particles often require the least or shortest treatment in the kiln but because of their physical size are subjected to the longest relative treating times.

The throughput velocities in processes of the type described have generally been set to accommodate the shortest residence time materials in order to provide for suflicient'time to permit adequate treatment of these short residence time materials. It is apparent, that in this situation, the longer residence time materials, that is the smaller particles, are subjected to too long treatment times. In addition, since the treatment is keyed to the short residence times of the larger particles, the throughput of any given kiln is reduced as compared to what it would be for any given particle size material except the largest. I i

Processes of this general type are used for magnetizing roasting of iron ore, for direct reduction of ironand/or nickel ore to form sponge iron or sponge nickel respectively, for the solubilization of phosphates with the aid of sodium carbonate slag or relatively. pure sodium carbonate, or for the treatment of zinc containing residues among other processes. Some of these processes have special problems peculiar to them which have not-been particularly met and solved by the prior art treating process as generally described above. For example, in the direct reduction of fine grained iron ore to form sponge iron, the kiln charge passes through a state in which the constituent particles tend to agglomerate even when such particles are below the melting temperature of either the ore being treated or the softening temperature of the treating material where such is a solid. It has been found that this state is generally reached when reduction of the iron ore has proceeded to an extent suflicient to have provided about 20 to 50 percent by weight metal in the kiln charge. Where this agglomeration occurs, the agglomerating particles tend to occlude other solids which are existent in the kiln charge, such as for example, solid treating materials such as coal. Additionally, the charge particles which are in an agglomerating state, are subject to and tend to adhere to the kiln walls whereby they represent a loss in terms of products and further represent an efficiency loss by reason of their tendency to reduce the available kiln volume and therefore reduce the available kiln throughput.

It has been proposed to attempt to remedy the agglomeration and related problems by using larger than required quantities of treating material solids, e.g. coal, in order to disperse and dilute the metal ore in the state of reduction. It had been thought that by keeping the metal particles being reduced in a more dispersed phase, it would be possible to prevent or at least minimize their tendency to agglomerate. This may or may not accompliSh the intended purpose however, great detriments result from its use regardless of whether it succeeds in minimizing the tendency for the ore charge to agglomerate. The greatest disadvantage, resultant from the use of excess treating material, e.g. coal, is economic in nature. The surplus coal or other solid treating agent is not readily recoverable from the product and therefore remains a part thereof to the economic loss of the process. This excess coal could possibly be disposed of during the process by burning; in which case there is not only the economic loss occasioned by the cost of the excess coal but also the utility cost necessary to furnish the additional air or oxygen necessary for combustion of the excess coal.

There are also technological disadvantages. Where excess coal is used, additional heat must be added to the kiln in order to heat the greater mass of material required for a given feed of ore and recovery of product. This heat input can be maintained at the same level as would have been required for the same ore charge and product recovery without regard for the excess coal or other treating material, but this would necessitate operating the kiln at a lower overall temperature and thus re tion-goes down; the throughput of a given kiln must necessarily sufier.

One method, which has received some attention, for minimizing the above recitedproblems associated with the use of widely varying ore feed to reducing kilns has been to reduce the variation in particle size of theme feed. This has been accomplished by abstracting the finest particle sizes from the ore feed and pelletizing them whereby larger particles are produced which are more in conformity with the large sizes of particulate ore being fed. To produce pellets from small particle ore stock, it is most desirable and possibly even essential, that the particles being pelletized have a content of about 80 weight percent of particles having a size below about 60 microns. This is equivalent to about 1400 to 1800 square centimeters per gramJWhere the particles to be pelle'tized have'less than this quantity of less than 60 micron particles; it-is necessary to grind or in other manner reduce the particle size to conform to this standard. Thus, there is economic disadvantage to this pelletizing process not only in the pelletizing operation costs themselves, but also in any associated size reduction costs which may be incurred.

Unpelletized, fine particle size ore is generally charged directly as such to a reducing kiln only in those cases where a subsequent operation is to be carried out, such as melting in an electric furnace, and where the reduction required of the charged ore, need be carried only to about 70 percent of completion.

Since much of the available ore and other treatable materials have particle sizes ranging from the fine particle size of about 0.09 to 1.0 millimeter to the coarser particle sizes of about 2 to millimeters, it would be desirable to provide a process for reduction or other treatment of various particle size material without the necessity of subjecting the charge material to preconditioning or sorting steps or operations.

It is therefore an object of this invention to provide a novel process for the treatment of particulate solids.

It is another object of this invention to provide a process for reducing varying size particulate ore;

It is a further object to provide a novel process-for reducing iron and/or nickel ores to their metallic-constituents in a more efficient manner than has been available in the prior art.

Other and additional objects of this invention will become apparent from a consideration of this entire specification including the claims appended hereto.

In accord with and fulfilling these objects, this invention resides, in one of its aspects, in a process for treating particulate solids having a distribution of particle sizes, wherein said particles being charged are separated, as by screening, for example, into three (3) grades; below about 60 microns; between about 60 microns and about 5 millimeters; and above about 5 millimeters. :The fraction between about 60 microns and about 5 millimeters is fed through the lower end (discharge end) of otherwise conventionally operating kiln. The varying particle size charge stock is propelled into the kiln in such a' manner that the particles travel a distance Within the'kiln substantially proportional to their size. That is, the large size particles tend to travel within the kiln a distance which is proportional to their size and relatively greater than the distance travelled within the kiln of the smaller particle sized ore.

Thus, it is practical, to control the residence time of various particle sizes of ore within the kiln in direct proportion to their size rather than in inverse proportion as was the case in the prior art. The larger particles have a longer residence time by reason of having been propelled into the kiln a greater distance and having to travel a greater distance in the kiln to reach the kiln exit. The smaller particles on the other hand proceed into the kiln a shorter distance, thus have a shorter distance to travel in the kiln than the larger particles. It is thereby practical, to properly control the residence time of" each particle according to its size with large particles being afforded a longer residence time than smaller particles.

It is preferred in the practice of this invention that the charged particles distribute themselves over at least about 4 meters of kiln length since it has been found that for the particle size range set forth, this distance is sufficient to provide for a quantitative distribution of particles according to their size assuming that all of the particles have about the same density.

It the ore to be treated in the kiln has a fraction having a particle size larger than about 5 millimeters, such fraction is preferably separated and fed to the kiln through or near the upper end (charge end) thereof. The fines of the ore to be treated, that is that fraction of the ore having a particle size less than microns, are suitable pelletized by known techniques to agglomerate them into larger particles.

Pellets having a particle size above 5 millimeters are preferably fed to the kiln from the upper end. Pellets having a particle size less than 5 millimeters are preferably charged to the kiln by propelling from the lower end of the kiln.

It is within the spirit and scope of this invention to propel the entire original ore to the kiln without a preliminary separation should it be found that there are only small amounts of fines below 60 microns and only small amounts of large particles above 5 millimeters. This modified operation may introduce minor inefiiciencies into the process however, it is believed that such minor inefiiciencies will be more than compensated for by the economic advantage of not requiring a separate preliminary grading appartus and process or a separate feeding device at the upper end of the kiln.

It may be practical in some modes of operation, particularly where substantiallyall, or most, of the charge is propelled into the kiln only to about half or quarter of the'kiln length, to provide throttling means at the kiln exit so as to provide in the kiln for any desired residence time of feed particles. In this same manner, it is practical to provide for whatever filling degree of material in the kiln is desired in any particular operation.

When the process according to the invention is applied to the direct reduction of iron ore to metallic iron it is suitable to maintainthe contents of the kiln at a temperature above 800 C., preferably 9001000 C. at leastin the range in which the relatively fine-grained iron ore which is blown in falls into the charge. The process according to'the invention may be generally carried out with a co-current or counter-current flow. Where a counter-current flow isused, the finer. particles being blown in at the discharge end, theusual advantages of a counter-current flow operation are obtained together with the further advantage that the distribution of the various particle size classes over the length of the kiln is particularly well in accordance with the residence time required for each of these classes. 7 e

The carrier gas for the solids to be blown in includes suitably at least part of the combustion air which is introduced through the central burner.

When a more uniform distribution of the injected solids over a particularly large part of the length of the kiln bed is desired, it may be suitable to blow the solids through a plurality of pneumatic conduits which protrude into the interior of thekiln. These conduits are then suitably arranged at dilferent inclinations relative to the kiln axis and/or are operated at different gas velocities and/or difierent gas pressures. For the same purpose it may be desirable to .supply the injecting carrier gas under a pulsating pressure.

It may be considered desirable or practical to preheat the ore being charged to the kiln. This may be carried out in any manner considered suitable. One practical method of accomplishing this preheating is to provide a heat exchange system using the kiln exhaust gases as the heating medium. These exhaust gases may be blown over or through the ore to be kiln-treated. It is preferred to preheat the ore to a temperature up to the kiln reaction temperature, since by this procedure the required residence time in the kiln is shortened. It is within the scope of this invention to utilize the preheat gases, or at least a portion thereof, as the conveyor gases by means of which the ore to be treated is blown into the kiln. If these preheat gases are used as the pneumatic conveyor for the charging of ore, it is required that these gases not contain substantial amount of material which is detrimental to the kiln operation. It is of course preferred that the pneumatic conveying gases contain nothing which is detrimental to the kiln operation.

If insufficient heat sources are economically available to preheat the feed ore to a temperature up to the kiln operating temperature, it may be practical to preheat the ore to a somewhat lower temperature, e.g. about 300 to 900 C. rather than the preferred 900 to 1100 C. It also may be practical to supplement heat exchange type heating, where such is insufliciently economically available, with direct fired type of preheating. This can oftenbe accomplished by providing appropriate conditions to afterburn exhaust gases normally used for preheating by heat exchange alone. If this supplemental direct fired heating is employed, it is often necessary to add oxidizing agent, suitably air or oxygen, to these exhaust gases in order to support combustion thereof. This after-burning can be accomplished during or prior to heating of the feed ore as desired.

Where the heat exchange preheating gases, either with or without after-burning, are used to heat the feed ore in the form of passing these gases through the particles of ore to be kiln-treated, it may be practical to combine this step with the size grading of the feed ore. This can be accomplished in the known manner by utilizing the preheating gases as the size sorting medium.

The fine fraction and the fraction with a particle size between about'60 microns and about millimeters of the ore to be processed in the kiln is preferably at least partially fed to the kiln through feeders suitably spatially located within the kiln. Shell burner or air inlet tubes have been found to be eminently well suited to use for this purpose. These devices are conventionally known in the art. Such shell burners and/or air inlet tubes are described in, for example, U.S.- Pats. "2,829,842 and 3,029,141.

It has been found practical to introduce at least part of the combustion supporting material, e.g. air, into' the kiln through the samemeans as the fine fraction of the ore is fed through. In this manner, the total amount 'of extra equipment associated with and attached to the kiln is kept to a minimum. It is also practical to provide, in the same manner as the fine fraction of the ore, for the introduction of suitable solid reducing agent such as coke or coal. Such reducing agent may suitable be introduced simultaneously with but separately from the ore to'be treated. It is preferred however, to charge the kiln with an appropriately proportioned mixture of reducing agent and ore with both in equivalent particle sizes such that each is available to the other in the kiln.

It is within the scope of this invention to utilize as a feed ore, material which contains sulfide or carbonate. In fact such material can be treated advantageously by the process of this invention by first feeding this ore' to roasting operations, suitably fluidized bed reactors or multiple hearth kiln, and then feeding the roasted product to the kiln of this invention, preferably a rotary kiln, wherein it is reduced as above described.

Thisprocess can also be used to advantage in processing sulfur containing ores, particularly iron and/or nickel ore. In the processing of these types of ores, it is the usual practice to add a sulfur combining mixture to the ore. Conventional materials known for this purpose are lime and dolomite among others. In so treating such ore, it is known, that the sulfur combining admixture adheres quite tenaciously to the solid reducing agent, e.g. coal etc., used. It has been found to be difficult, if not impossible, to remove the deposited sulfur combining material from the particulate reducing agent, particularly when the product discharged from the kiln, is subjected to wet classification. It is usual to recirculate excess reducing agent in the kiln product back to the kiln in order to make greatest economic use of this material. When such coated reducing agent is recirculated back to the kiln, hydrate of lime and/or magnesium hydroxide are introduced into the kiln. Conventionally these materials react with the ore to produce iron compounds which are themselves extremely difficult to reduce and which adversely effect the reducibility of the entire ore charge. It has been found that this marked economic disadvantage is largely overcome by the practice of this invention, if this coated coal or other reducing agent is introduced into the kiln at the discharge end thereof. Thus, the coated coal comes into contact only with substantially reduced metallic material which has a much lower tendency to react therewith to form compounds which show a great resistance to reduction.

A further aspect of this invention, particularly applicable to such situation, where a sulfur combining agent is required in the reduction of iron and/or nickel ores, is to adjust the relative particle sizes of the sulfur combining agent and the reducing agent such that the sulfur combining agent is present in smaller particle sizes than is the reducing agent. This practice severely reduces the tendency of the sulfur combining agent to deposit on and coat the surplus reducing agent. The kiln product may then be separated by dry methods, e.g. conventional screening, thereby further reducing the tendency of the sulfur combining material to tenaciously adhere to the surplus reducing agent discharged from the kiln. The relatively large particles of recovered coal to be recycled may be fed through the top of the kiln or at any other point convenient and/or according to this invention. The fine particle size reducing materials on the other hand, which still have a tendency to pick up coatings of the sulfur combining .materials can be fed in the vicinity of the kiln product discharge as set forth above.

Ithas"been found practical to use sulfur combining materials in particle sizes of about 0.1 to 1 millimeter and reducing agent in particle sizes of about 1 to 3 millimeters. The fine particle size reducing agent recycle referred to above, suitably has a particle size below about '1 millimeter.

Inaddition to the reducing agents generally referred to herein above, it is also within the spirit and scope of this invention to use substantially any carbonaceous solid fuel such as in addition to coke and anthracite coal, lignite, peat, brown coal', bituminous coal or other similar products. These reducing agents may be used singly or in admixture with each other or with other reducing agents. These materials may be utilized in the rough form as delivered or they may be ground to the size preferred by the operator as desired. It is also possible to pelletize these solid carbonaceous materials either alone or in admixture as desired.

EXAMPLES The following examples were carried out in a rotary kiln of 7.8 meter length and an interior diameter of 0.5 meter. The rotary kiln was equipped with shell burners and a central burner through which gas and/or air was delivered into the kiln to control the temperature over the kiln length. The central burner was designed in such a manner that it was not only possible to deliver gas or air but also to feed pneumatically coal and fine grained ore or a mixture of said materials. The central burner was mounted by a sealing device at the discharge end of the kiln in such a manner that it was possible to vary the inclination of the burner with respect to the kiln axis.

Example 1: Magnetising roasting of iron ores A low grade iron ore containing 37% Fe total was roasted with lignite-coal as reducing agent i.e. the ore was reduced from hematite to magnetite. This material was ground and concentrated by magnetic separation.

The ore was ground before feeding into the kiln. The grain size distribution was as follows:

Grain size, mm. Grain size distribution, percent The ground ore was separated in two fractions by screening the fraction with a grain size +3 millimeters was fed to the kiln in the conventional manner from the upper end (charge ends). The fraction with a grain size -3 millimeters was mixed with coal and pneumatically charged through the lower end (discharge end) of the kiln. The carrier medium was air which was preheated by the waste gas of the kiln. The reduction coal contained:

Percent Ashes 10.6 Fix carbon 49.7 Volatile constituents 37.6

The above values being based on dry material. The net calorific value was 5790 kcaL/kg. The grain size was 2 millimeters.

The feeding rate was:

160 kg./h. ore +3 mm. 140 kg./h. ore 3 mm.

20 kg./h. coal 2mm.

The temperature in the reduction zone was regulated at 800 C. by introducing gas through the central burner.

The reduced ore was water quenched, ground and concentrated by known multiple wet magnetic separation. The concentrate contained 68% Fe total. The yield of Fe was 90%.

Example 2: Prereduction of iron ores }mixture For producing pig iron in an electric furnace fine grained ores were prereduced with high volatile coal in a rotary kiln and hot charged to the electric furnace.

The iron ore contained:

Percent Fe total 55.7 TiO 14.1 A1 0 4.6 SiO 1.7 CaO 0.1 MgO 1.3 S 0.1

The grain size distribution was:

Grain size, mm. Grain size distribution, percent +18 5.53

The net calorific value was 6710 kcaL/kg.

The ore was separated in 2 fractions by screening. The

8 fraction with a grain size +6 mm. was fed to the kiln from the upper end in a ratio of 135 kg./h. The fraction 6 mm. was mixed with coal 1() mm. and pneumatically charged into the kiln through the lower end. The carrier medium was combustion gas with a temperature of 600 C. The pneumatically charged material was distributed over a length of 5 meters on the surface of the charge in the kiln, whereby the coarse particles were carried farther into the kiln than the finer particles.

The pneumatically charged mixture consisted of Kg./h. Ore 45 Coal 60 The temperature in the reduction zone was 1050- 1100", C. The temperature was regulated on a constant value by feeding air through the shell burners causing combustion of the volatile matter in the kiln gases.

The reduction product was reduced to an amount of 49%. It contained:

Percent Fe total 51.3 Fe metallic 14.0 Fe++ 302 Fixed carbon 12.3

The material was hot charged into an electric furnace.

'Example 3: Reduction of specular iron ore concentrate produced on Humphrey spiral concentrator.

The grain size distribution'of the ore was:

Grain size, mm. Grain size distribution, percent The ore contained ca. 66% Fe total in the form of hematite, The reducing agent was a high volatile pit coal with the following composition:

Percent H O 7.5 Ashes 7.2 Fixed carbon 56.6 Volatile constituents 28.2

The net calorific value was 6540 kcaL/kg. The cakingcapacity (Damm method) was about 7.5. The swellingindex was in the range from 2-2.5. The coal was used in a grain size -5 mm.

The ore was separated by air separation in a fraction +0.09 mm. was pelletized with hot heavy oil on a pelletizing disk. The pellets had a diameter of 1 to 3 mm. and were pneumatically charged into the kiln together with the fraction +0.09 mm.

The feeding rate was:

- Kg./h. Spiral concentrate 140 Raw coal -5 mm. 160 Return coal 5 mm. Dedusted dolomite 1 mm. 7.2

Return coal and dolomite were charged in known manher at the upper end of the kiln. Spiral concentrate (fraction +0.09 mm. and pellets) were mixed with raw coal and charged through the lower end of the kiln by using preheated combustion air (500 C.) as carrier medium.

The reduction product was cooled to a temperature below C. in a volating cooling cylinder which was cooled indirectly.

The cooled material was separated in a magnetic and a non-magnetic fraction by magnetic separation. The

magnetic fraction was reduced greater than 95% and contained 90.5% Fe total and 87% Fe metallic.

Example 4. Prereduction of nickeliferous ores =For production of ferro-nickel in an electric low pit smelting furnace lateritic nickel ore was prereduced with lignite and heavy oil in a rotary .kiln. The ore had the following composition:

The ore was separated into a fraction +2 mm. and a fraction 2 mm. The fraction 2 mm. was mixed with 7.5% lignite-coal (from the same composition as that of Example 1) and during sprinkling with heavy oil agglomerated.

The fraction +2 mm. was charged to the upper end of the rotary kiln. The mixture of agglomerated ore and coal was pneumatically fed through the lower end. The reduction temperature was about 1020 C. and was regulated by additional combustion of oil and burning the volatile constituents of coal and oil.

.The reduction product contained:

Percent Ni 2.6 Fe 80.0

That means that the nickel content was metallized to an amount above 80% and the iron content to about 25%. The content of remaining fixed carbon was about 0.3%.

The prereduced ore was hot charged into the smelting furnace.

What is claimed is:

1. In the process for reducing iron oxide containing ore to sponge iron in a treatment zone of a rotary kiln, which kiln has a discharge end from which treated solid material and in which kiln endothermic reactions involving solids havng difierent particle sizes are carried out; which process comprises feeding said ore and a solid carbona- 10 ceous reducing agent to said treatment zone; maintaining the temperature and pressure in said treatment zone sufficient to sustain said treatment; and recovering the solid product resulting from such treatment at said discharge end, the improvement which comprises separating said iron oxide containing ore to be treated into a first fraction comprising particles under 60 microns, a second fraction comprising particles 60 microns to 5 mm. and a third fraction comprising particles over '5 mm. spatially distributing the particles of said second fraction in said kiln by pneumatcally propelling such into said kiln from the discharge end of said kiln into the treating zone of said kiln whereby the particles of said second fraction spatially distribute themselves in said zone over at least 4 meters of the length of said zone, wherein said penumatic propulsion propels the larger particles of said second fraction more distantly from the discharge end of said kiln than the smaller particles of said second fraction, thereby providing a residence time in said zone of said fine grained ore particles substantially proportional to the size of such particles; forming said first fraction into pellets of at least about 5 mm. and feeding such pellets and third fraction comprising coarse particles of the iron oxide containing ore of at least 5 mm. size to the feeding end of said zone.

2. The improved process claimed in claim 1, wherein said second fraction of said ore is preheated by exhaust gases from said treatment zone.

3. The improved process claimed in claim 1, wherein said iron oxide containing ore contains nickel and said treatment zone operates at about 800 to 1100 C.

References Cited UNITED STATES PATENTS 1,171,360 2/1916 Schwahn 26 X 1,490,012 4/ 1924 Kapteyn 7526 1,524,182 1/1925 Kjolberg 7526 1,736,665 11/1929 Pape 7526 2,075,823 4/1937 Mullen et a1. 7526 X 3,126,275 3/1964 Tudja 7533 X 3,206,299 9/1965 Senior et a1. 7536 X 2,823,108 2/ 1958 'Gerlach 7533 3,034,884 5/1962 Meyer et al. 1,717,160 6/1929 Kichline 7531 2,994,601 8/1961 Greene 7531 3,138,451 6/1964 =Gerlach 7533 3,224,871 12/1965 Collin 7534 3,238,039 3/1966 Sasabe 7533 X HENRY W. TAR-RING, Primary Examiner US. Cl. X.R. 7582 (mum) m PATENT m me?) (ZEEJQEEQEFlCATE (PE? COwElREUllUN Patent NO; 3,667,933 Dated June 6, 1972 Inven tor (g0 GUENTER HEITMANN It is certified that error appears in the above-identified patent and that said Letters Patent are hereby corrected as shown below:

Col. 1, line 30, delete the comma after "particles" 7 C01. 1, lineS38-34, "deposiited" shouldbe deposited-- Col. 5, line 57, "suitable" should be "suitably-- Col, 9, line 52, after "material" insert --is discharged and a feeding end into which ore is charged,--

Signed and sealed this 1mm day Of November 1972.

(SEAL) Attest:

EDWARD M.FLETCHER,JR. ROBERT GOTTSCHALK Attesting Officer Commissioner of Patents ORM PC4050 H069) USCOMM-DC 0037mm;

U S, GOVIRNMENT PRINTING OFFICE I959 0"16633 

